Copper Grade Research Articles

Влияние структуры и состава технически чистой бескислородной меди различных производителей на ее электропроводность и механические свойства

The composition, structure, electrical conductivity and mechanical properties of oxygen-free copper grade C10100 with a purity of 99.997% produced by Sichuan Huazi Technology Co., Ltd (China) intended for use in the production of niobium-titanium superconductors as a stabilizing component of a multi-filament composite were studied. It was shown that the recrystallization onset temperature of deformed copper samples in the form of wire is about 150 °C, and as a result of collective recrystallization at temperatures above 300 °C, the grain size reaches a value of more than 180 μm. The RRR parameter characterizing the electrical conductivity of copper at the cryogenic temperature of samples in a fully recrystallized state reaches a value of 1300 units, which is three times higher than the level typical for copper grade C10100 produced by leading manufacturers of cryogenic copper for superconductors. Maximum plasticity (up to 50%) of copper wire samples is achieved after short annealing (less than 5 minutes) at a temperature of 300°C. Increasing the annealing temperature to 600°C leads to a decrease in mechanical properties due to grain growth and the well-known effect of temperature plasticity failure.

Assessing the role of seawater in the flotation behavior of chalcopyrite and sphalerite

ABSTRACT In seawater flotation, some chemical properties of seawater affect the flotation separation of chalcopyrite and sphalerite. Using seawater instead of fresh water to separate copper – zinc can save fresh water resources and reduce lime consumption, which has broad prospects in terms of economic and environmental protection. In this paper, the flotation separation behavior of chalcopyrite and sphalerite in seawater was tested and the mechanism was analyzed. As a contrast, the copper grade of concentrate in seawater flotation declined from 23.7% to 22.6%. Meanwhile, the recovery of chalcopyrite declined from 66.2% to 42.8% and the SI value declined from 2.210 to 1.599. With the addition of sodium hexametaphosphate, sodium silicate and EDTA, the SI value increased from 1.599 to 2.176, 2.238 and 2.096, respectively. Nevertheless, when EDTA was excessive, the SI value was only 1.998. This indicated that excessive EDTA is unfavorable to flotation separation of chalcopyrite from sphalerite. Wettability analysis, zeta potential analysis, X-ray photoelectron spectroscopy, Fourier transform infrared spectroscopy, density functional theory calculation and E-DLVO calculation verified the flotation separation behavior and revealed the mechanism.

Flotation of Copper Sulfide Ore Using Ultra-Low Dosage of Combined Collectors

Copper sulfide ores frequently co-occur with pyrite, presenting a significant challenge for their selective separation during beneficiation processes. Despite advancements in flotation technology, there remains a critical need for efficient methods to enhance copper recovery while suppressing pyrite interference, particularly without compromising the associated precious metals such as gold and silver. Current practices often struggle with achieving high selectivity and recovery while maintaining environmental sustainability. Here, we investigate the efficacy of a ternary collector mixture consisting of ammonium dibutyl dithiophosphate (ADD), butyl xanthate (BX), and ethyl xanthate (EX) for the selective flotation of copper sulfide from a complex ore containing 0.79% Cu and associated precious metals (0.233 g/t Au and 5.83 g/t Ag). A combination of lime and hydrogen peroxide as inhibitors was employed to suppress pyrite effectively under alkaline conditions (pH = 11.33). The results demonstrate that the optimized ternary collector system (ADD:BX:EX at a ratio of 1:0.5:0.5) significantly improves the copper grade and recovery at an ultra-low dosage of 10 g/t. The optimized flotation method using the combined collectors and inhibitors effectively separated chalcopyrite from pyrite, achieving a copper concentrate with 20.08% Cu content and a recovery of 87.73%. Additionally, the process yielded notable recoveries of gold (9.22%) and silver (26.66%). These findings advance the field by providing a viable and environmentally conscious approach to the beneficiation of sulfide ores, potentially serving as a blueprint for processing similar mineral deposits while minimizing reagent usage and costs.

Statistical Evaluation of Flotation Behavior of Chalcopyrite in the Presence of SIPX and Acetoacetic Acid n-Octyl Ester as a Novel Collector Blend: A Sustainable Approach

The chalcopyrite deposit in Malanjkhand, India, is the largest copper ore producer in the country. However, its flotation performance has room for improvement. This study used standard flotation experiments using a mechanical flotation cell using a conventional collecting reagent and a collector blend consisting of xanthate and ester. A three-factor, three-level Box-Behnken design was used to statistically evaluate the experimental design. The obtained data were analyzed using an ANOVA, cubic plots, and response surface methods. The goal was to evaluate the effects and interactions of three key process parameters in chalcopyrite flotation: the dosages of sodium silicate (depressant), sodium isopropyl xanthate (collector), and acetoacetic acid n-Octyl ester (co-collector/modifier). The results implied that introducing acetoacetic acid n-Octyl ester along with minor tweaks in the dosages of sodium isopropyl xanthate helped increase copper grades by at least 15%, with good recovery percentiles. Among the three parameters tested, the copper’s grade and recovery were considerably positively influenced by the AoE dosage in the collector blend. Employing 0.006 kg/t sodium silicate and 0.0065 kg/t sodium isopropyl xanthate with 0.005 kg/t of acetoacetic acid n-Octyl ester, an optimum copper recovery of 88.87% could be achieved. However, with a sodium silicate dosage of 0.0048 kg/t, a SIPX dosage of 0.008 kg/t, and an AoE dosage of 0.005 kg/t, optimized copper grade (1.55%) could be achieved, which is a 78.1% increase from the feed sample grade. To validate the expected results, verification experiments were carried out, and the experimental findings were found to be on at par with the statistical model predictions. Furthermore, the calculated SPI value of 0.0000147cap Kg−1 for a global index per resident lies between 0 and 1.

Bio-mediated enhancement of supergene copper mineralization: Evidence from Cu isotope geochemistry

The relationship between microbial activity and the supergene modification of ore systems has been a major subject of debate. Here, we present isotopic evidence of microbial-driven secondary copper mineralization in the active cementation zone of the Las Cruces deposit, a volcanogenic massive sulfide deposit located in Iberian Pyrite Belt, Spain. Copper isotopic data show that the lower isotopic ratios (δ65Cu ≈ −9.2 ± 0.11 ‰, the lowest value measured worldwide in a supergene environment) are found in the upper part of the cementation zone, the same zone where the maximum copper grades are found and where there is direct evidence of extremophilic microbial activity. There is a tendency towards higher values downwards through the cementation zone (−9.2 ± 0.11 ‰ to + 1.67 ± 0.11 ‰ δ65Cu) and upwards into the former Cu-depleted gossan that originally capped the cementation zone (−7.79 ± 0.11 ‰ to −1.32 ± 0.11 ‰ δ65Cu). As microbes preferentially sequester the lighter isotope when incorporating intracellular Cu, this distribution indicates that microbes played a major role in the formation of the high-grade zones. Water arrives to the deposit enriched in isotopically heavy copper, likely because it has leached other ore bodies upstream. δ65Cu values of water currently flowing into the system are remarkably more positive than those in the ore, indicating that microbial activity is a major cause of copper isotope fractionation. At least half of the copper transported by the incoming waters remains within the ore body. Our best interpretation is that the large and high-grade cementation zone at Las Cruces is of biogenic origin, and that the primary mineralization acted as a trap for copper transported by groundwater, leading to the formation of an exotic mineralization distal to sub-eroded massive sulfides located upstream.

Replacive IOCG systems in the Ossa Morena Zone (SW Iberia): The role of pre-existing ironstones as a geochemical trap

The central Ossa Morena Zone (SW Iberia) hosts a regionally extensive ironstone level interbedded with bimodal volcanic rocks, limestone and shale of Lower-Middle Cambrian age. The stratabound ironstone includes dominant magnetite and hematite with locally abundant chert and barite. It is interpreted as being (sub-)exhalative at or near the seafloor and formed during a rifting event that postdated the Cadomian orogeny. In some places, such as in the Las Herrerías deposit, the ironstone is irregularly replaced by a chalcopyrite-rich ore; the Cu-rich mineralization is accompanied by the pervasive phyllic alteration of the hosting siliciclastic sediments. The highest copper grades are found when the ironstone is crosscut by WNW-ESE-trending late-Variscan extensional brittle-ductile structures that are interpreted as the feeder channels for deep hydrothermal fluids. A similar nearby copper-rich mineralization (Pallares) is is likely controlled by the tectonic contact between limestone and pyrite-rich black shale.Sr-Nd whole-rock isotope geochemistry data suggests that the Sr in the ironstone (87Sr/86Sri ≈ 0.7088) is close to isotopic equilibrium with the local exhalative barite (0.7084–0.7086) and Cambrian seawater. The ironstone has a significantly more crustal εNd initial signature (<-1.8) than the coeval volcanic rocks (+5.2 to + 7.9). The younger sulfide mineralization inherited the Nd isotope composition of the ironstone but shows a significant enrichment in 87Sr (87Sr/86Sr > 0.7091) that is interpreted as related with the input of genetically different and more crustally-derived hydrothermal fluids.39Ar-40Ar dating of the phyllic alteration suggest that the copper mineralization was formed at ca. 332–330 My. These ages are coeval with those of small peraluminous granite intrusions that host Cu-Au vein-like mineralization and dated at 331.8 ± 1.6 Ma (LA ICPMS U-Pb zircon). Our interpretation is that the copper-rich mineralization at the Las Herrerías area is the distal expression of an intrusion-related hydrothermal system.Numerical modelling shows that ironstone is an effective trap for copper precipitation due to the large changes in pH and fO2 that take place when copper-bearing acid and reduced fluids react with the brittle ironstone. The precipitation of chalcopyrite, however, is controlled by the amount of available reduced sulfur in the ore trap. The δ34S values of the sulfides (+12.6 to + 21.6 ‰) suggest that the most likely source for the reduced sulfur is the thermogenic reduction of aqueous sulfate equilibrated with the exhalative barite (δ34S, +31.4 to + 35 ‰) with some minor input of reduced sulfur leached from the metasediments.This system could be considered as a variant of the IOCG clan. The formation of the ironstone and the copper mineralization, however, are separated by more than 200 My. Probably, many IOCG systems have a similar origin as Las Herrerías, with an ironstone being just a passive geochemical trap for the copper–gold mineralization.

Effect of APMImBr ionic liquid on the migration of silver in copper electrorefining

As the grade of copper concentrate decreases and its composition becomes increasingly complex, the silver content in the anode plate casting after fire refining increases. This leads to a high silver content in the copper cathode during electrorefining and a substantial loss of precious metals. This study investigated the effect of 1-aminopropyl-3-methylimidazolium bromide (APMImBr) ionic liquid on the migration of silver during high-silver anode plate electrorefining. The results show that APMImBr is adsorbed on the anode surface by reacting with Ag+ released from the anode to form AgBr precipitates, thereby preventing Ag+ ions from entering the electrolyte. Simultaneously, agglomerates of fine Ag–Se compounds and AgBr particles form, resulting in anode slime particles that are sufficiently large to settle quickly to the bottom of the cell, thus inhibiting fine silver-containing particles from adhering to the cathode. Furthermore, APMImBr adsorbs onto the cathode surface, forming CuBr intermediates and inhibiting Cu2+ electrodeposition. This facilitates the refinement of the surface grains and flattens the copper cathode surface. The Ag content of the cathode decreased continuously as the concentration of APMImBr increased. The Ag content of the cathode decreased by 41.67% from 6.0 ppm (no additives) to 3.5 ppm upon the addition of the optimal concentration of 300 mg/L of APMImBr.

Optimising the placement of additional drill holes to enhanced mineral resource classification: a case study on a porphyry copper deposit

ABSTRACT In this study, we compared outcomes of optimising the placement of five additional drill holes using three geostatistical cost functions (AKV, WAKV, and CV) and the Particle Swarm Optimisation algorithm (PSO). WAKV identified locations with higher average copper grades compared to AKV. Conversely, CV suggested sites with high kriging variance and copper grade variation. Initial holes, alongside those determined by each cost function, were used to classify mineral resources. Findings underscored the effectiveness of optimising drill hole placement based on cost functions in reducing uncertainty and improving mineral resource classification.

An eco-friendly depressant derived from grape seed extract (GSE) as depressant for flotation separation of pyrite from chalcopyrite under low-alkalinity conditions

This study explores the selective depression mechanism of pyrite by proanthocyanidins (PC) during chalcopyrite-pyrite flotation separation. The selective inhibition effect of inhibitor PC was examined through flotation experiments. The depression mechanism of PC on pyrite was analyzed employing X-ray photoelectron spectroscopy (XPS) analysis, selective adsorption tests, adsorption experiments, contact angle measurement, and zeta potential measurement. The artificial mixed mineral flotation test of chalcopyrite-pyrite at a ratio of 1:12 was conducted under pH 7, PC concentration of 75 mg/L, and PAX concentration of 18.75 mg/L. The copper concentrate exhibited a recovery rate of 88.97%, with a copper grade of 29.43%. The sulfur grade in copper concentrate was 40.35%, and the pyrite recovery rate was 1.97%. PC can be used as a potential depressant of pyrite. adsorption experiments and Zeta potential measurement indicated that PC treatment significantly impeded PAX adsorption on the pyrite surface, as confirmed by XPS test results and Ion adsorption analysis.

ImhoflotTM Flotation Cell Performance in Mini-Pilot and Industrial Scales on the Acacia Copper Ore

The present work investigates a comparative study between mechanical and ImhoflotTM cells on a mini-pilot scale and the applicability of one self-aspirated H-16 cell (hybrid ImhoflotTM cell) on an industrial scale on-site. The VM-04 cell (vertical feed to the separator vessel with 400 mm diameter) was fabricated, developed, and examined. The copper flotation experiments were conducted under similar volumetric conditions for both the ImhoflotTM and mechanical flotation cells keeping the rest of the parameters constant. Further, one H-16 cell was positioned at four different stages in the Gökirmak copper flotation circuit of the Acacia (Türkiye) copper beneficiation plant, i.e., at (i) pre-rougher flotation, (ii) rougher concentrate, (iii) cleaner-scavenger tailing, and (iv) first cleaning concentrate aiming at enhancing the flotation circuit capacity through flash flotation in the rougher stage, reducing copper grade in the final tailing, and increasing cleaning throughput, respectively. Comparative copper flotation tests showed that ultimate recoveries using the ImhoflotTM and mechanically agitated conventional cells were 94% and 74%, respectively. The industrial scale test results indicated that locating one pneumatic H-16 cell with the duty of pre-floating (also known as flash flotation) led to the enrichment ratio and recovery of 4.84 and 89%, respectively. Positioning the H-16 cell at the cleaner-scavenger tailings could diminish the copper tailings grade from 0.43% to 0.31%. Further, a relatively greater enrichment ratio and copper recovery were obtained using only one ImhoflotTM cell (1.76 and 64%) in comparison with employing four existing mechanical cells (50 m3, each cell) in series (1.45 and 60%) at the first cleaner stage.

Influence of Mineralogy and Mineralogy Approach to Optimize Processing: A Case Study of Tin–Copper Polymetallic Ore

Tin-Copper polymetallic ore is a type of typical ore that cassiterite is closely associated with sulfide minerals. In mineral processing of tin–copper polymetallic ore, flotation is generally used to recover valuable sulfide minerals, while gravity separation is used to recover cassiterite. A mine in Yunnan, China, uses the traditional “flotation–gravity separation” process to recover copper and tin but faces several problems during processing, such as an insufficient copper grade in Cu concentrate, a much higher grade of As in S concentrate, and a grade of S in Sn concentrate that exceeds the standard. A process mineralogy study was conducted, with a focus on Cu–S mixed concentrate, S concentrate, and Sn rough concentrate. It was determined that the main cause of these problems is not the liberation or size distribution of valuable minerals but the superstructure of pyrrhotite, which represents one of the most abundant minerals in the products. Based on EMPA, SEM-EDS, and XRD data, both monoclinic pyrrhotite and hexagonal pyrrhotite occurred in all samples. The abundance of different superstructures of pyrrhotite in one sample was determined by means of particle extraction and area calculation from microscopic images, and the distribution characteristics of monoclinic pyrrhotite and hexagonal pyrrhotite in the whole process were clarified. This process mineralogy study indicates that the strong magnetic hexagonal pyrrhotite mainly affects the copper recovery during flotation, and the hexagonal pyrrhotite mainly affects the recovery of cassiterite during gravity separation. Strong magnetic monoclinic pyrrhotite and weak magnetic hexagonal pyrrhotite should be fully considered in the optimization of mineral processing, and the magnetic separation of pyrrhotite should be adopted to optimize the overall environment of copper flotation and tin gravity separation.

Open-Pit Pushback Optimization by a Parallel Genetic Algorithm

Determining the design of pushbacks in an open-pit mine is a key part of optimizing the economic value of the mining project and the operational feasibility of the mine. This problem requires balancing pushbacks that have good geometric properties to ensure the smooth operation of the mining equipment and so that the scheduling of extraction maximizes the economic value by providing early access to the rich parts of the deposit. However, because of the challenging nature of the problem, practical approaches for finding the best pushbacks strongly depend on the expert criteria to ensure good operational properties. This paper introduces the Advanced Geometrically Constrained Production Scheduling Problem to account for operational space constraints, modeled as truncated cones of extraction. To find the best solution for this problem, we present a parallel genetic algorithm based on a genotype–phenotype model such that the genotype symbolizes the base block of a truncated cone, and the phenotype represents the cone itself. A central computer node evaluates these solutions, collaborating with various secondary nodes that evolve a population of feasible solutions. The PGA’s efficacy was validated using comprehensive test instances from established research. The PGA solution exhibited a consistent average copper grade across periods, with its incremental phases reflecting real-world mine geometry. Moreover, the benefits of the MeanShift clustering technique were evident, suggesting effective phase-based scheduling. The PGA’s approach ensures optimal resource utilization and offers insights into potential avenues for further model enhancements and fine-tuning.

Extraction of Copper from Pregnant Leach Solution (PLS) and Reduction of Crud Formation

ABSTRACT Presented universal recommendations for the reduction of crud formation, reviewing various aspects affecting the loss of diluent and extractant due to the formation of a third phase (crud), encountered in many hydrometallurgical plants using copper solvent extraction. In this paper introduce the reasons of the crud formation , description of the stable emulsions containing solid particles and choice of technological scheme. An analysis of operations to be considered in the operation of copper solvent extraction plants is presented, due to the deterioration of copper grade below 0.5% in ore and the increase in aluminum content up to 20–25%. Results of use of extraction properties of modified extractants of ACCORGA series and unmodified extractant Lix984N on pregnant leached solution of the most important deposits (Almaly and Aktogay) are given. It is established that for solutions of Almaly deposit, ACCORGA5640 and unmodified Lix984 extractants have high selectivity to copper. Distribution of iron, silica, and copper ions during extraction was studied, and it was found that high selectivity to copper/iron and copper/silica is possessed by extractant Lix 984N, ACCORGA5640. It was found that for solutions of the Aktogay deposit, the ratio of interphase suspension in both extractants is the same, and phase separation in experiments with the use of extractant Lix 984N was 10 s shorter with the use of extractant ACCORGA5640. It was found that the addition of ACCORGACR60 reagent in the amount of 5-10 ppm also leads to a decrease in the volume of crud by 13%, for a stronger suppression of crud.

Relation of bubble size, grade and recovery in the copper flotation systems

In recent years, studies have shown that bubble structure, which is affected by reagent types, dosages, and particle size, is closely related to flotation performance. In contrast to two-phase systems, there is very limited information regarding bubble size in three-phase systems. To this aim, we investigated the relationship between bubble size and flotation performance with two different copper ores. Firstly, the surface tension and bubble size measurements were carried out in two-phase systems with Methyl isobutyl carbinol, and Polypropylene glycol methyl ether. Then the bubble size measurements were performed in three-phase systems at laboratory and industrial scales. The results showed that surface tension decreased with higher frother amounts and higher pHs. Although remarkable decreases in bubble size were observed with increasing frother amount, a direct relationship between surface tension and bubble size was not observed. However, pH affected the behavior of the frother, resulting in smaller bubble sizes at pH 12 compared to pH 7. Similarly, the three-phase laboratory tests showed that higher amount of frother caused a reduction in the bubble size and the copper grade. Increasing the amount of collector also caused the bubble size and copper grade to decrease, but the frothers were more effective in bubble size reduction. Compared to the two-phase, in the three-phase study bubble size was notably bigger due to the presence of a collector and solid. In the plant scale, smaller bubble size and higher grade were obtained compared with the laboratory scale due to presence of cleaning circuit measurements. The froth phase properties were also affected by the ore types due to properties such as size and hydrophobicity of the various mineral species present.

Study on harmless treatment and valuable elements recovery of neutralization residue from acid wastewater of copper mine

Recently, the common practice for treating metallurgical acidic wastewater involves neutralization with lime emulsion, resulting in residue that is typically landfilled or stockpiled. This contributes to long-term land occupation and environmental pollution. In this study, neutralization residue from a mine underwent comprehensive analysis using XRD, XRF, SEM-EDS, and other techniques. Utilizing acidic wastewater from the local mine as a leaching agent, the leaching characteristics of the residue were systematically investigated, and subsequent toxicity testing was conducted. Copper and iron, enriched in the leachate, were recovered employing the “pre-iron removal – sulfurization – pH compliance” process. Leaching experiments revealed the feasibility of mitigating toxicity by adjusting the pH of local mine wastewater (pH 1.5) with sulfuric acid and reacting with the residue at a ratio of 3:1 (mL/g), under conditions of 25 °C for 120 min. Under these optimized conditions, the recovery rates for Cu, Fe, As, Zn, and Ni reached 99.42%, 96.71%, 88.89%, 99.23%, and 98.21%, respectively. Toxicity testing classified the leaching residue as non-hazardous general solid waste for discharge. Recovery experiments demonstrated optimal conditions for the process, including pre-iron removal at pH 4, H2O2 10 mL/L 30 min, 25 °C, sulfurization with NaHS at 2 g/L, 20 min, 25 °C, and an emission process at pH 7.5, 60 min and 25 °C. The pre-iron removal residue (grade 45.54%) is conducive to entry into the iron plants for refinement processes aimed at enhancing its grade before entering the market, while the sulfide residue, with a copper grade of 57%, hold commercial potential. The remaining liquid adhere to established discharge standards. This process is characterized by a waste-to-waste and waste-to-treasure paradigm, demonstrating robust process stability.

A bulk sorting trial of copper ore using a magnetic resonance analyser

Sensor-based bulk ore sorting (BOS) is a technique that may lead to uplift of processing grades by excluding waste as well as an increase in usable resource by sorting material designated as mineralised waste. A magnetic resonance analyser (MRA) was installed on the primary crushed ore feed to the sulphide plant at the Kansanshi copper mine in northwestern Zambia to trial BOS. The conveyor belt operated at 2800tph and carried ore of which greater than 85% by weight of the copper-bearing minerals are chalcopyrite. The MRA was therefore tuned for chalcopyrite’s zero static field magnetic resonance. The analyser achieved a 3σ detection limit of 0.045 wt% copper as chalcopyrite over a 4 s measurement time and standard belt loading, with lower detection thresholds possible if the measurement time increased.A trial of ore diversion was then carried out at the site. Pods of ore measured as less than 0.2 wt% copper as chalcopyrite by the MRA were diverted by a dead box diverter system to form a low-grade stockpile of approximately 10,000 tonnes. A cavity in the mill feed stockpile was arranged and the rejected material fed back through the crusher, past the MRA and directly through the milling circuit to the flotation feed where a Vezin sampler was used to take a sample every 5 min. This sample was then tested for copper grade by the site metallurgical laboratory. The MRA measured a grade of 0.16 wt% copper as chalcopyrite over the test time period and the Vezin samples for the last hour of the trial reported approximately 0.2–0.3 wt% copper despite some mixing with unsorted ore being inevitable. This is as clear a demonstration of bulk ore sorting (BOS) as practically achieveable in the production environment at Kansanshi and, to the author’s knowledge, the first demonstration of BOS on a copper ore stream of at least this throughput.

Predictive insights for copper recovery: A synergistic approach integrating variability data and machine learning in the geometallurgical study of the Tizert deposit, Morocco

This work involved the application of machine learning (ML) to predict copper recovery through variability data in the context of the geometallurgical study of the Tizert deposit in Morocco. For this purpose, geological and metallurgical variability investigations were conducted on drilling samples from Tizert ore, leading to the consolidation of a robust database for ML algorithms. The dataset comprised 345 observations across seven variables, namely GPS coordinates, geological units, lithological facies, the main copper deportment, and metallurgical responses from flotation tests, as well as the copper grade and the oxidation rate. The geological and mineralogical variability revealed two geological units: The Basal Series and Tamjout Dolomite unit, hosting copper mineralization dominated by chalcocite and malachite. Besides, the metallurgical evaluation demonstrated recovery variability, primarily influenced by oxidized copper. Copper oxides exhibit less efficient flotation than sulfides, with recovery decreasing as the oxidation rate increases. The prediction of copper recovery was performed using the variability data through the XGBoost algorithm. The resulting model showed good performance on the test dataset, with a coefficient of determination R2 of 0.83 and a RMSE of 2.47, improving the accuracy of copper recovery predictions. In addition, the model elucidated the most significant variables impacting Tizert ore behavior, notably the casing, copper-bearing mineral, and oxidation rate. These variables were considered key factors for advancing Tizert ore valuation and geometallurgical model development. Overall, metallurgical modeling based on the ML approach provided an innovative way to integrate variables into mine planning and processing. The use of such a predictive model gave valuable insights for improving efficiency, understanding the metallurgical behavior, optimizing production, and reducing costs along the value chain of the Tizert deposit. Further data processing efforts are underway, to improve the assessment of copper recovery quality for the Tizert ore and the development of a complete geometallurgical model on a plant scale.

Effect of different process water sources on rougher flotation efficiency of a copper ore: A case study at Sarcheshmeh Copper Complex (Iran)

In this research, the effect of different sources of process water on the flotation efficiency of copper sulfide ore prepared from the Sarcheshmeh copper mine was investigated. For this purpose, samples of fresh water to the plant, overflows of copper-molybdenum concentrate thickener, copper concentrate thickener, and recycled water pool as well as a mixture of fresh water and recycled water were prepared and characterized. Flotation tests were performed under the same conditions as the plant’s rougher circuit and were kept constant during all experiments. Grade and recovery of copper, iron, molybdenum, and silica were selected as the metallurgical response of flotation tests. The results were subjected to statistical analysis to assess the relative significance of which water source affects the flotation performance as evaluated from the experimental results. The results showed that the copper concentrate thickener overflow had the greatest effect on the flotation efficiency, so the grade and recovery decreased by about 10% and 75% for copper, and 10% and 6% for iron in the concentrate, respectively, while the grade and recovery increased up to 0.1% and 12% for silica, and 3% and 25% for molybdenum, respectively. The reason for this effect was attributed to the high content of suspended solid particles, and Cu2+, Mo2+, and Fe2+ cations in this water source that increased the coating effect over gangue minerals and entrainment rate. The improvement of molybdenum flotation was also ascribed to the possible presence of residual diesel oil from the flotation process in the plant. Due to the relatively equal amount in all sources of process water, the effect of anions and ions of dissolved salts was difficult.

Tourmaline breccia pipes of the Giant Copper porphyry system: Extending the Cascadia porphyry district into southern British Columbia, Canada

ABSTRACT The A. M. breccia is part of the Giant Copper property of British Columbia, Canada. It is the only well-defined tourmaline breccia pipe (TBP) in the Canadian Cordillera. The A. M. breccia shares similarities with other TBPs, most notably those of South America. The A. M. breccia demonstrates concentric zonation with regard to breccia texture, consisting of an outer rim of shingle breccia surrounding a fragmental breccia core. Copper grades correlate to breccia zonation, with higher Cu grades within shingle breccias relative to the fragmental core. Hypersaline fluid inclusions were identified in quartz cement within the A. M. breccia, which reflects conditions expected in porphyry copper deposits. Breccia textures, mineralization, and fluid inclusions at the A. M. breccia are nearly identical to other well-studied porphyry-related TBPs. The strong similarities suggest the A. M. breccia is a porphyry-related TBP. This interpretation encourages deep drilling to identify further mineralization within the A. M. breccia pipe as well as discovery for conventional porphyry mineralization at depth. Similarities can also be drawn to Ancestral Cascadia arc porphyry copper deposits and associated TBPs. The proximity of Giant Copper to the Ancestral Cascadia porphyry district infers a genetic relationship and the interpretation that the Giant Copper porphyry system developed from magmatism related to the Ancestral Cascadia subduction. Classifying Giant Copper as an Ancestral Cascadia porphyry system extends the district into southern British Columbia, where Giant Copper marks its northernmost extent.

Reclamation of iron and copper from BCL slag in Botswana

High-grade copper ores have been depleted over the years, making it a challenge in the mining industry. This investigation focused on a methodology to recover iron (Fe) and copper (Cu) from a copper/nickel slag obtained from the Bamangwato Concession Limited (BCL) mine in Botswana. In this modified flotation approach, the Response Surface Methodology (RSM) was used in conjunction with the Central Composite Design (CCD) and Analysis of Variance (ANOVA) to obtain the best optimal flotation conditions for the recovery of iron and copper. Using the RSM – CCD methodology, the optimal predicted responses were illustrated by a coefficient of determination R2 = 0. 9839 for recovery for Fe and 0.9655 for recovery for Cu. The recovery of copper increased with the increasing dosage of Na2S and collector dosage, while the increase of pH, had a decrease in recovery of copper due to the decline in the stability of the froth, which led to the resistance to form stable bubbles for efficient recovery of copper. Selective flotation of copper and iron was achieved by varying the Na2S dosage to achieve maximum recovery. Under these flotation conditions of PAX (800 g/t), pH (8), −75 μm, sulfurizing agent (Na2S, 1000 g/t), flotation time of 8 min, pH regulator of NaOH and H2SO4 and Methyl Isobutyl Carbinol (MIBC) from the experimental runs merited a grade upgrade of Cu in froth concentrate from 0.581 mass% to 0.884 mass%. An enrichment ratio of 2 was realised, with the recovery of Cu being 62%, whereas Fe in the froth concentrate increased from 69.8 mass% to 71.8 mass%. The main aim was to upgrade the grade and recovery of copper and iron to enhance the recovery for copper and iron in the next experimental stage of leaching.